Process for Leaching Lateritic Ore at Atmospheric Pressure

ABSTRACT

An atmospheric leaching process in the recovery of nickel and cobalt from a lateritic ore, the lateritic ore including a low magnesium ore fraction and a high magnesium ore fraction, the process including the steps of: (a) forming an aqueous pulp of the lateritic ore, (b) leaching the aqueous pulp with a concentrated mineral acid at atmospheric pressure to produce a slutty containing a pregnant leach liquor and a leach residue, (c) treating the pregnant leach liquor either separately or as part of the slurry to recover dissolved nickel and cobalt therefrom, leaving a magnesium barren solution, (d) treating the magnesium containing solution to recover a magnesium containing salt therefrom.

This application is a continuation of and claims priority to PCTapplication PCT/AU2005/001497 filed Sep. 30, 2005, published in Englishas WO 2007/035978 on Apr. 5, 2007, the entire contents of each areincorporated herein by reference.

The present invention relates to a hydrometallurgical process to recovernickel and cobalt from nickeliferous laterite ores and, in particular,to the atmospheric leaching of both low magnesium fraction (e.g.limonite) and high magnesium fraction (e.g. saprolite) ores with amineral acid to dissolve nickel and cobalt. In a preferred form of theprocess, the process also includes a step in which magnesium values inthe leach liquor are recovered.

The known reserves of nickel and cobalt in nickeliferous oxide ores,e.g. those referred to as laterites comprising limonite and saproliteore, are far greater than the corresponding reserves in sulfide ores. Animportant disadvantage when processing laterite ores, however, is theinability to beneficiate these ores by conventional techniques.

A number of new hydrometallurgical processes are being developed for theextraction of nickel and cobalt from nickeliferous laterite ores. Manyof these processes require the dissolution of the metal values withsulfuric acid at high temperature (245°-270° C.) and pressure (525-785psig), followed by solid-liquid separation and neutralization ofresidual free acid present at ambient pressure. This is the basic “MoaBay Process”, as described by J. R. Boldt and P. E. Queneau in “TheWinning of Nickel”, Methuen, London, 1967. In this process, thenickeliferous ore is first made into a pulp having a solids content ofabout 40% before leaching at high temperature and pressure. Duringpressure leaching most metals dissolve and iron and aluminum arerejected by hydrolysis to hematite and alunite, respectively. Afterleaching, the pulp is cooled and washed by counter current decantationand the solids are directed to tailing treatment. Excess acid isneutralized and the remaining iron and aluminum are precipitated ashydroxides with the addition of alkali. Nickel and cobalt aresubsequently recovered via sulfide precipitation.

Several variations of the high-pressure acid leach (HPAL) method havebeen devised with the aim of improving the process and economicalaspects. For example, U.S. Pat. No. 4,044,096 provides guidelines tooptimize the high-pressure acid leaching of nickeliferous lateritic oresthrough a combination of operational steps to improve the economics andefficiency of leaching. The steps include scalping laterite ore toremove the coarse (high magnesium) fraction and thus lower the acidconsumption.

The HPAL process is most amenable for high iron ores containing 40 wt %iron or higher. Lateritic ores with an iron content less than 40 wt %contain in general a higher amount of acid consuming minerals and aretherefore not preferred for direct high pressure leaching. U.S. Pat. No.3,804,613 teaches a method of high-pressure acid leaching of saproliteore at relatively low acid/ore ratios by preconditioning the saprolitewith leach liquor from the high-pressure leach step. No mention is madeof concurrent limonite leaching.

U.S. Pat. No. 3,991,159 teaches the use of saprolite ore to neutralizeacid resulting from the high-pressure acid leach of limonite ore.Leaching of the saprolite fraction is carried out at high temperature(150°-250° C.) and pressure for effective iron and aluminum rejection,but with relatively low nickel extraction from the saprolite ore. Inanother process, U.S. Pat. No. 4,097,575 teaches saprolite ore roastingat 500°-750° C. under oxidizing conditions to increase itsneutralization capacity before neutralization of HPAL liquors. Thisprocess suffers from the additional need for roasting facilities.

While the prior art HPAL methods obtain a high extraction of nickel andcobalt, they require the use of expensive equipment and sophisticatedmaterials of construction to withstand the use of concentrated acid atthe high temperatures needed (200°-300° C.). Several alternatives to theHPAL process to recover nickel and cobalt from laterite ore have beenproposed.

For example, U.S. Pat. No. 4,062,924 describes a method for leachinglimonite ores in acidic media at temperatures up to 110° C. and in thepresence of hydrogen sulfide gas to precipitate dissolved nickel andcobalt. Most dissolved iron is also reduced to the divalent oxidationstate however, consuming very high amounts of the reducing gas inaddition to high acid consumption. U.S. Pat. No. 4,065,542 teaches asimilar method. In this process, ferrous iron produced by the methoddescribed above is used to leach metal values from manganiferous seanodules. U.S. Pat. No. 4,511,540 illustrates a way to recover nickel andcobalt from ores with a manganiferous matrix by leaching with sulfuricacid in the presence of sulfur dioxide gas at temperatures below theboiling point of the liquid solution. None of these processes includesthe treatment of saprolitic ores.

In the process of U.S. Pat. No. 3,793,432, limonite ore is leached withsulfuric acid at a pH below 1.5, while simultaneously adding alkalineiron-precipitating agents. The process is carried out at atmosphericpressures, but requires leaching times in excess of 40 hours and usuallyfrom 60 to 100 hours for efficient nickel extraction and ironprecipitation. No use of saprolite is made in this process. U.S. Pat.No. 4,410,498 teaches a method to leach saprolite ore with sulfuric acidat atmospheric pressure, while adding a reducing agent to maintain theredox potential between 400 and 600 mV. In another process, described inU.S. Pat. No. 5,571,308, nickel and cobalt are leached from saproliteore by contact with a mineral acid at room temperature or in thetemperature range of 60°-80° C. The leaching mode can be conducted byheap, vat, or agitation leaching.

U.S. Pat. No. 6,261,527 also discloses a hydrometallurgical process forthe recovery of nickel and cobalt from both limonite and saprolite ores,however in that process, iron is rejected as jarosite.

There are environmental concerns with this iron removal process as thejarosite compounds are thermodynamically unstable. Jarosite maydecompose slowly to iron hydroxides releasing sulphuric acid. Thereleased acid may redissolve traces of precipitated heavy metals, suchas Mn, Ni, Co, Cu and Zn, present in the leach residue tailing, therebymobilizing these metals into the ground or surface water around thetailings deposit.

Another disadvantage of this process is that jarosite contains sulphate,and this increases the acid requirement for leaching significantly.Sulphuric acid is a large input in acid leaching processing, so there isalso an economic disadvantage in the jarosite process.

The present invention aims to overcome or alleviate one or more of theproblems associated with prior art processes. The discussion ofdocuments, acts, materials, devices, articles and the like is includedin this specification solely for the purpose of providing a context forthe present invention. It is not suggested or represented that any orall of these matters formed part of the prior art base or were commongeneral knowledge in the field relevant to the present invention as itexisted before the priority date of each claim of this application.

SUMMARY

According to the present invention, there is provided an atmosphericleaching process in the recovery of nickel and cobalt from a lateriticore, said lateritic ore including a low magnesium ore fraction and ahigh magnesium ore fraction, said process including the steps of:

-   (a) forming an aqueous pulp of said lateritic ore,-   (b) leaching said aqueous pulp with a concentrated mineral acid at    atmospheric pressure to produce a slurry containing a pregnant leach    liquor and a leach residue,-   (c) treating the pregnant leach liquor either separately or as part    of said slurry to recover dissolved nickel and cobalt therefrom,    leaving a magnesium containing barren solution,-   (d) treating said magnesium containing solution to recover a    magnesium containing salt therefrom.

An advantage of the invention is the provision of an efficient andeconomical method to leach both low magnesium (e.g. limonite) and highmagnesium (e.g. saprolite) ores in a single process stage at atmosphericpressure, to obtain high percent dissolution of nickel and cobalt. Afurther advantage of the method is that it avoids the high capital costsassociated with sophisticated autoclaves. Another advantage of apreferred form of the method is that it also avoids the production ofjarosite. An advantage of a preferred form of the invention, is that themagnesium containing barren solution produced from the leaching processis treated to recover magnesium sulphate, which is then processed togive MgO, Mg(OH)₂ or MgCO₃ and SO₂. The SO₂ is advantageously used toregenerate H₂SO₄. The MgO or MgCO₃ may be fed back into the leachingprocess as a neutralising agent, disposed of as a stable residue, orsold as a commercial product.

BRIEF DESCRIPTION OF THE DRAWINGS

FIG. 1 is a graphical depiction of calcination of magnesium sulphateheptahydrate in a thermogravimetric analyser (TGA) under a flow of drynitrogen over time.

FIG. 2 is a graphical depiction of calcination of magnesium sulphateheptahydrate in a thermogravimetric analyser (TGA) under a flow of dryhydrogen over time.

FIG. 3 is a flow sheet of one embodiment of the present invention inwhich the recovery of Ni and Co is carried out using one of mixedhydroxide precipitation, mixed sulphide precipitation, solventextraction, or ion exchange.

FIG. 4 is a flow sheet of another embodiment of the present invention inwhich Ni and Co are recovered using a resin-in-pulp extraction techniqueprior to the removal of residual iron and aluminium and with subsequentmanganese precipitation and separation of leach residue from the barrensolution.

DETAILED DESCRIPTION OF THE INVENTION

The low magnesium containing ore fraction includes the limonite fractionof the laterite ore (Mg wt % approximately less than 6). This fractionmay also include low to medium level magnesium content smectite ornontronite ores which generally have a magnesium content of about 4 wt %to 8 wt %. The high magnesium containing ore fraction includes thesaprolite fraction of the laterite ore (Mg wt % greater thanapproximately 8). This fraction may also include smectite or nontroniteores. The formation in step (a) of an aqueous pulp of both the lowmagnesium and high magnesium containing ore fractions is generallycarried out in sodium, alkali metal and ammonium free water at solidsconcentration from approximately 20 wt % and above, limited by slurryrheology.

The ratio of acid to combined ore is typically at least 0.5. Preferably,the ratio is about 0.5 to 1.0, such as 0.5 to 0.7.

The aqueous pulp is subjected to a leaching step in step (b) utilising aconcentrated mineral acid at atmospheric pressure. Preferably leachingis conducted whilst agitating leach reactants. Typically the leachingstep is carried out at a temperature up to the boiling point of theleach reactants at atmospheric pressure. Most preferably the reactiontemperature is as high as possible to achieve rapid leaching atatmospheric pressure. A preferred leaching temperature is at least 60°C., more preferably at least 75° C. In a preferred embodiment, leachingis carried out at around 80° C. or higher, such as at least 85° C. Inanother preferred embodiment, leaching is conducted at around 95° C.

Preferably, the leaching of both the low and high magnesium fractionsoccurs in a single process stage, which may comprise a single step, inwhich the two fractions are leached simultaneously, e.g. in the sametank or reactor. Alternatively, the two fractions may be leached insequential steps in the single process stage. In that case, preferablythe low magnesium fraction (e.g. limonite) is leached in a first step,and then the higher magnesium fraction is subsequently added to theslurry to be leached in a second step. The sequential leaching of thelow and high magnesium fractions may be in accordance with thedisclosure of WO 03/093517, the entire disclosure of which isincorporated herein by reference.

Leaching is conducted for a period of time sufficient to release atleast a substantial portion of the nickel and cobalt from the lateriteore into solution. Typically leaching is conducted for up to 30 hours.However, preferably leaching is conducted for up to 5 hours. Morepreferably, leaching is conducted for up to 4 hours. In a preferredembodiment, leaching is conducted for about 2 hours.

The leaching process typically also results in precipitation in at leastsome of the iron in the ore as one or more Fe containing compounds, suchas a sulphate, a hydroxide or an oxide.

The mineral acid used in the leaching process is preferably sulphuricacid, more preferably it is concentrated sulphuric acid. Theconcentration of sulphuric acid added to the ore pulp is preferablygreater than 90 wt %. The dose of sulphuric acid is preferably 100 to140% of the stoichiometric amount required to dissolve approximatelyover 90% of nickel, cobalt, iron, manganese and over 80% of the aluminumand magnesium in the ore.

The ratio of the high magnesium ore to low magnesium ore is ideally in adry ratio of from 0.5 to 1.3. Preferably, the ratio is from 1 to 1.30.However, the high/low magnesium ore ratio will largely depend on thelaterite ore composition.

The leaching of both the high and low magnesium fractions may optionallybe followed by a second leaching step. In the second leaching step, anyunused acid from the first leaching step may be reacted with additionalhigh magnesium ore fraction, such as saprolite. Leaching conditions oftemperature, time and acid concentration are typically similar to thoseof the first leaching step.

Addition of saprolite can cause further precipitation of Fe containingcompounds.

Conditions of temperature, time and acid concentration may convenientlybe controlled to allow part or all of the iron and aluminum to beprecipitated. The acidity may be conveniently controlled by the additionof saprolite, MgO, Mg(OH)₂, MgCO₃ or another alkali. For example, theleach slurry may be treated in accordance with the method disclosed inWO 03/093517 (the entire disclosure of which is incorporated herein byreference), in which saprolite ore is added to a leach slurry in orderto precipitate goethite or other relatively low sulphate-containingforms of iron oxide or iron hydroxide. Alternatively, the leach slurrymay be treated in accordance with the method disclosed in U.S. Pat. No.6,261,527 (the entire disclosure of which is also incorporated herein byreference) in which an iron precipitating agent selected from sodium,potassium ammonium ions and mixtures thereof is added to the slurry toprecipitate jarosite.

In a preferred embodiment, MgO is added to the slurry in order toprecipitate iron containing compounds. Preferably, the MgO additionresults in an increase of pH to a value of 3.0 or higher, causing ironprecipitation.

The leached slurry is then treated to recover dissolved nickel andcobalt values therefrom. Such metal extraction treatment may be one ormore of techniques known to those working in the art. Examples of suchmetal extraction techniques include ion exchange, resin-in-pulp, directrecovery by solvent extraction, mixed hydroxide precipitation or mixedsulphide precipitation. Preferably the recovered nickel and cobaltvalues are recovered as mixed nickel/cobalt hydroxides or mixednickel/cobalt sulphides.

Prior to or during recovery of nickel and cobalt from the leach liquor,the solid leach residue which usually includes precipitated ironcompounds such as Fe sulphates e.g. jarosite or Fe hydroxides, e.g.,goethite, may be removed from solution depending on the recovery processused. Alternatively, the solid residue may be retained with the leachsolution during subsequent removal of residual Fe and/or Al.

Prior to or after recovery of nickel and cobalt from the leach solution,the spent leach solution is preferably treated to remove any residual Feand/or Al in solution. Typically, this step requires an increase insolution pH, such as by adding a neutralising agent, such as MgO,Mg(OH)₂ or MgCO₃, and preferably addition of an oxidising agent such asair. Typically a sufficient quantity of neutralising agent is added suchthat the solution pH is increased to around 3 or above. A sufficientamount of the oxidising agent is also added to oxidise any residual Fe²⁺in solution to Fe³⁺, which then precipitates out as goethite.

After removal of Ni, Co, Fe and Al from the spent leach liquor, thesupernatant solution mainly contains dissolved magnesium, possiblytogether with a small quantity of manganese. The supernatant solution isthen treated in order to recover the magnesium as magnesium salts. Thisis achieved typically by evaporation until the magnesium saltscrystalise out. Alternatively, reverse osmosis or precipitation by astrong alkali such as caustic soda, soda ash or lime, may be used.

The magnesium salt is typically a magnesium sulphate where the leachingacid used was sulphuric acid. It has been the conventional practice todiscard the magnesium salts as waste, meaning that metal values in thesalts are therefore lost. Moreover, when the magnesium salt comprisesmagnesium sulphate, the sulphate component is also lost, which increasesthe acid requirement for the leaching process significantly. Sulphuricacid is usually an expensive input in acid leaching, so there is aneconomic disadvantage in simply discarding a source of sulphate.

Accordingly, in a preferred embodiment of the process, the presentinvention is also concerned with treating the magnesium salt to recovermagnesium compounds. Where the magnesium salt comprises MgSO₄, therecovery process also preferably includes a sulphate recovery stage.Preferably, the magnesium is recovered as a magnesium oxide, magnesiumhydroxide or magnesium carbonate. More preferably, the magnesium isrecovered as magnesium oxide. The magnesium recovery process maycomprise that disclosed in co-pending Australian provisional patentapplication 2005900431 filed on 1 Feb. 2005, the entire disclosure ofwhich is incorporated herein by reference. Alternatively, the magnesiumsalt may be subjected to calcination. Where the magnesium salt ismagnesium sulphate, calcination results in formation of MgO and/or MgCO₃and SO₂ gas. The SO₂ gas may be captured and fed to a sulphuric acidproduction process, in which sulphuric acid is regenerated according tothe following process:

SO₂+H₂O+½O₂=H₂SO₄

The MgO, Mg(OH)₂ or MgCO₃ produced from the magnesium salt is a goodsource of alkaline compound, which can be fed back to the leach solutionas a neutralising agent to effect precipitation, separately or incombination, metals such as Ni, Co, Al, Fe, Mn and other elements asdesired.

EXAMPLE 1

A mixture of limonite and saprolite ore in a dry ratio of about 1 isformed into an aqueous pulp. The aqueous pulp is then mixed withconcentrated sulphuric acid, having a concentration of 93% H₂SO₄, toform a leach slurry. The dose of acid is greater than 100% of thestoichiometric amount required to dissolve over 90% of the Ni and Co inthe combined ore fractions. A first leaching process is conducted in asingle reactor at a temperature of at least 80° C. and for at least 2hours. During the first leaching process, iron compounds precipitate outof solution.

Overflow from the leaching process is conveyed to a second reactor,where a saprolite ore slurry is added to the mixture. A second leachingprocess is then conducted, also at a temperature of at least 80° C. andfor a time of around 2 hours. During the second leaching process,further iron compounds precipitate out from solution.

After completion of the second leaching process, the solid residue isseparated from the leached slurry. The pregnant leach solution is thensubjected to a recovery process during which nickel and cobalt valuesare recovered.

The spent leach solution is also treated to remove any residual iron andaluminium. This is effected by the addition of a neutralising agentcomprising MgO or MgCO₃. The pH of the barren solution is therebyincreased, to a value higher than 3. The iron is precipitated largely ashydroxides, such as Fe(OH)₃.

At this time, the barren leach solution contains mainly dissolvedmagnesium. The spent leach solution is directed to an evaporation pondand excess water evaporated therefrom, causing crystallisation ofmagnesium sulphate.

The magnesium sulphate is then subjected to a magnesium recoveryprocess. This comprises calcination to produce MgO, or MgCO₃, and SO₂gas. The SO₂ gas is then used as a reactant in a sulphuric acid recoveryprocess.

The following two Examples are concerned with the recovery of magnesiumfrom magnesium sulphate crystals. Example 2 is a Comparative Exampledemonstrating calcination of MgSO₄.7H₂O under non-reducing conditions,which shows that MgSO₄ remains as the product. However, Example 3demonstrates that calcination under reducing conditions achievesproduction of MgO at moderate temperatures, that is, at temperaturessignificantly lower than those at which calcination is conventionallyconducted.

EXAMPLE 2—COMPARATIVE EXAMPLE

A sample of magnesium sulphate heptahydrate (4.0353 g) was placed in asmall crucible and calcined in a thermogravimetric analyser (TGA) undera flow of dry nitrogen (5 L/min). The temperature in the TGA was raisedby 10° C./min from room temperature to 1000° C. The sample exhibited aweight loss of approximately 2.07 g by 400° C. and exhibited very littlefurther weight loss. The resulting mass of the sample (1.9386 g)corresponds closely with the formula MgSO₄ (theoretical weight of 1.9706g). A graphical depiction of the TGA run is shown in FIG. 1.

EXAMPLE 3

A sample of magnesium sulphate heptahydrate (4.0093 g) was placed in asmall basket and calcined in the thermogravimetric analyser (TGA) undera flow of dry hydrogen (5 L/min). The temperature in the TGA was againraised by 10° C./min from room temperature to 1000° C. The sampleexhibited a weight loss of approximately 2.03 g by 350° C.,corresponding to the loss of waters of crystallisation. The weight thenremained stable until 630° C. at which a further weight loss ofapproximately 1.06 g. Rapid weight loss slowed at a temperature of 810°C. By the time 1000° C. had been reached the total weight loss wasapproximately 3.29 g. The remaining sample was carefully removed fromthe container and weighed. The mass of the weighed sample (0.63 g)corresponds closely with the formula MgO (theoretical weight 0.655 g). Agraphical depiction of this run is shown in FIG. 2.

EXAMPLE 4

Example 4 is a Flowsheet, illustrated in FIG. 3, which sets out theprocess stages in an embodiment of the present invention. In thisExample, separation of the leach residue from the pregnant leachsolution takes place prior to removal of residual iron and aluminium andrecovery of Ni and Co metal values. The recovery of Ni and Co iseffected using one of the techniques selected from mixed hydroxideprecipitation, mixed sulphide precipitation, solvent extraction or ionexchange.

EXAMPLE 5

Example 5 is a Flowsheet, illustrated in FIG. 4, setting out the processstages in a further embodiment of the present invention. In Example 5,the Ni and Co metal values are recovered, using the Resin-in-Pulp(R-I-P) extraction technique, prior to removal of residual iron andaluminum, subsequent manganese precipitation and separation of leachresidue from the barren solution.

The above description of the invention is illustrative of the preferredembodiments of the invention. Variations without departing from thespirit or ambit of the invention described herein are to be consideredto form part of the invention.

1. An atmospheric leaching process in the recovery of nickel and cobaltfrom a lateritic ore, said lateritic ore including a low magnesium orefraction and a high magnesium ore fraction, said process comprising: (a)forming an aqueous pulp of said lateritic ore; (b) leaching said aqueouspulp in a single process stage with concentrated sulphuric acid atatmospheric pressure to produce a slurry containing a pregnant leachliquor and a leach residue; (c) treating the pregnant leach liquoreither separately or as part of said slurry to recover dissolved nickeland cobalt therefrom, leaving a magnesium containing barren solution;(d) treating said magnesium containing barren solution to recover amagnesium sulphate therefrom; (e) treating said magnesium sulphate torecover SO₂ gas and a magnesium compound selected from MgO, Mg(OH)₂ andMgCO₃, said magnesium compound being recyclable as a neutralising agentfor treatment of at least one of the pregnant leach liquor and thebarren solution; and (f) forming sulphuric acid from said SO₂ gas forrecycling to said leaching step (b).
 2. The process of claim 1 furtherincluding the step of increasing the pH of at least one of the pregnantleach liquor and the barren solution to produce an iron containingprecipitate.
 3. The process of claim 1, wherein both the low and highmagnesium fractions are leached simultaneously during said leaching step(b).
 4. The process of claim 1, wherein the low and high magnesiumfractions are leached sequentially during said leaching step (b).
 5. Theprocess of claim 1, wherein the leaching step (b) is conducted at atemperature above about 60° C.
 6. The process of claim 5, wherein thetemperature is at least about 80° C.
 7. The process of claim 1, whereinthe ratio of high magnesium ore to low magnesium ore is in a dry ratioof from about 0.5 to about 1.3.
 8. The process of claim 2, wherein thestep of increasing pH further comprises adding a neutralising agent tothe at least one of the pregnant leach liquor and the barren solution toincrease the pH to a value of approximately 3 or above.
 9. The processof claim 2 wherein the step of increasing pH further comprises adding anoxidising agent in order to oxidise any residual Fe²⁺ to Fe³⁺ and causeits precipitation as a hydroxide.
 10. The process of claim 9, whereinthe oxidising agent is air.
 11. The process of claim 1, wherein thepregnant leach liquor is treated to recover at least one of nickel andcobalt using at least one of ion exchange, resin-in-pulp, directrecovery by solvent extraction, mixed hydroxide precipitation and mixedsulphide precipitation.
 12. The process of claim 11, wherein thepregnant leach liquor is treated to recover nickel and cobalt usingmixed hydroxide precipitation or mixed sulphide precipitation.
 13. Theprocess of claim 1, wherein the magnesium sulphate is treated to recovermagnesium.
 14. The process of claim 1, wherein the magnesium sulphate issubjected to calcination to produce SO₂ gas and at least one of MgO andMgCO₃.
 15. The process of claim 1, wherein the magnesium compound isused as a neutralising agent in step (c).
 16. The process of claim 1,wherein the magnesium compound is used to precipitate at least one ofiron, aluminium, and manganese from at least one of the pregnant leachliquor from (b) and the barren solution from (c).